FLOTATION OF ALUMINA FROM GIBBSITE BEARING–SHALE OF … · 2014-07-03 · However, the flotation of gibbsite by using amines as collectors can not be applied for the south western

Journal of Engineering Sciences, Assiut University, Vol. 36, No. 2, pp.443-460, March 2008

443

OPTIMIZATION OF ALUMINA FLOTATION FROM GIBBSITE BEARING-SHALE OF SOUTH WESTERN SINAI, EGYPT BY

USING SODIUM OLEATE AS A COLLECTOR

M. M. Ahmed a; G. A. Ibrahim b ; A. A. Elmowafy c and H. G. Ahmed d a ,b

Mining and Metallurgical Engineering Department, Faculty of Engineering,

Assiut University, Assiut 71516, Egypt c, d

Nuclear Materials Authority, Egypt

(Received December 12, 2007 Accepted January 8, 2008)

The effect of sodium oleate on the flotation of alumina was studied using samples of gibbsite bearing–shale of south western Sinai, Egypt. The

assays of the samples are 18.98% Al2O3, 15.45% SiO2, 12.83% Fe2O3,

14.87% CaO, 5.74% P2O5, 5.34% MnO, 0.86% K2O, 0.76% Na2O, 1.50%

trace elements and 23.65% loss on ignition. The aim of this research is to

upgrade alumina in the concentrate to be suitable for the industrial

applications. Various operating variables (i. e. pH, pulp density,

collector dosage, and particle size) affecting the flotation process of

gibbsite were studied. Sodium oleate used as an anionic collector; as well

as, sodium hydroxide was used as a pH modifier.

The optimum conditions obtained were as follows: pH = 11, collector

dosage = 2000 g/t, pulp density = 250 g/l and particle size = (-100+80)

µm. At these conditions, a concentrate having a grade of 40.1% alumina

with component recovery of 96.1% was obtained. The grades of the other

ore constituents in final concentrate were as follows: SiO2 was decreased

from 14.2% to 5.9%, CaO was decreased from 14.1% to 4.8%, Fe2O3 was

decreased from 13.1% to 5.4%, P2O5 was decreased from 4.3% to 3.2%,

MnO was decreased from 4.2% to 3.5%, while K2O was increased from

0.89% to 1.3%, Na2O was also increased from 0.69% to 1.2%, traces

were increased from 1.6% to 2.7%, and loss on ignition was increased

from 25.4% to 31.9%. The component recoveries of SiO2, CaO, Fe2O3,

P2O5, MnO, K2O, Na2O, and traces of the final concentrate were 21.4%,

17.5%, 21.2%, 38.4%, 42.9%, 75.3%, 89.6%, and 86.9%, respectively.

The mass recovery of final concentrate was 51.5%.

KEYWORDS: Gibbsite, alumina flotation, collectors, isoelectric point,

pH, pulp density, particle size

NOMENCLATURE

c assay of constituent in

concentrate, %

C mass of concentrate, gm

f assay of constituent in feed, %

F mass of feed, gm

Rc(c) component recovery of

constituent in concentrate, %

Rc(t) component recovery of

constituent in tailings, %

Rm(c) mass recovery of concentrate, %

Rm(t) mass recovery of tailings, %

t assay of constituent in tailings, %

T mass of tailings, gm

M. M. Ahmed et al.

444

1. INTRODUCTION

Gibbsite Al (OH) 3 is a crystalline aluminum trihydroxide which is used in the

manufacture of refractory products [1]. The most abundant structural element in the

earthۥs crust is aluminium. It composes eight per cent of the earth

ۥs mineable layer and

is most abundant as aluminium oxide or combined oxides in a wide variety of minerals

[2,3]. Due to their high corrosion resistance and mechanical strength to mass ratio,

aluminum alloys are used as major structural material in aircraft, buildings, machinery

parts, beverage cans and food warps [4,5]. In nature, aluminum occurs only in

combination with other elements and is a part of the crystal structure of many rock

forming minerals. Bauxite is the most important commercial ore of aluminum [5].

Bauxite consists of several hydrous aluminium oxide phases [boehmite (γAlO.OH),

gibbsite (Al (OH) 3), and diaspore (α AlO.OH) [2]. Besides of its main use as an

aluminum ore, bauxite is used also to manufacture refractory products, aluminous

chemicals, abrasives, and miscellaneous applications. These applications include

β−alumina solid state electrolytes and first–class refractories [4].

The major impurities of gibbsite of south western Sinai are quartz, hematite

and kaolinite. Other impurities such as, rare earths and alkali occur in small

percentages. Gravity, magnetic, or electrostatic separation methods are not suitable for

the concentration of gibbsite ore. This may be due to the very small significant

differences in the physical properties (density, magnetic and electrostatic susceptibility,

etc.) between the mineral phases (gibbsite, kaolinite and quartz), which are considered

the major constituents of the ore. Flotation process was then selected as a suitable and

effective method for the concentration and upgrading of alumina from gibbsite

bearing–shale of south western Sinai to make it suitable for the Bayer process.

Many researchers [4-13] have studied the flotation of alumina using different

types of collectors. Their results are summarized and explained below.

Flotation was selected as the most suitable method for the recovery of a high

purity gibbsite concentrate from the Brazilian bauxite ore using alkyl sulfates as

collectors. The results were good at pH value of 2, where the component recovery of

alumina was 97.4% and the grade of alumina was 93% at a particle size of (-177+3)

µm and a pulp density of 20% solids [4].

In other research, sodium oleate was used to float gibbsite from a mixture of

quartz, kaolinite, iron and titanium oxides (Guyana bauxite ore) at pH values between

10 and 11. At the optimum conditions a component recovery of more than 90%

gibbsite was obtained [6]. The maximum floatability of gibbsite, chamosite and

boehmite was achieved at a pH of 8 during the adsorption of laurylamine on all

minerals. The adsorption of laurylamine occurred at a pH value ranged between 4 and

8 [7].

However, the flotation of gibbsite by using amines as collectors can not be

applied for the south western Sinai gibbsite ore because a significant amount of quartz

is present. Bulut and Yurtsever [8] found that, sodium dodecyl sulfate, as an alkyl

sulfate collector, is effective in the acidic pH range during the flotation of alumina

from kyanite ore using different collectors. Dodecylamine hydrochloride (DAH), as a

cationic collector, is effective in the basic pH range and potassium oleate (KOL), as an

anionic collector, floats kyanite at pH value ranges from 6 to 9. Doss [9] reported that,

anionic surfactants such as the alkyl sulfates and alkyl sulfonates are adsorbed on

OPTIMIZATION OF ALUMINA FLOTATION FROM…… 445

alumina by means of electrostatic and hydrophobic bonding. Consequently, when the

alumina surface becomes negatively charged above pH 9, sulphonate adsorption

ceases, but at pH values below 9, sulphonate ions is adsorbed to the positive alumina

surface by electrostatic attraction.

Aluminum silicates such as pyrophyllite, illite and kaolinite are impurity

minerals in diasporic-bauxite, it being necessary to remove these silicate minerals to

improve the mass ratio of Al2O3/SiO2 for the Bayer process. Quaternary ammonium

salt (DTAL) was used to float diaspore at a pH values ranged from 6 to 7 [10]. The

surface charges of gibbsite particles were probed by potentiometric titration and

subsequently were analyzed to estimate intrinsic proton affinity constants of OH

surface groups. It was found that the point of zero charge could be achieved at a pH

value ranged from 8.1 to 9.6 [11]. The electron spin resonance (ESR) was used to

study the adsorption of spin probe analogs of stearic acid on amorphous alumina [12].

The point of zero charge (ZPC) of gibbsite is at a pH value of 9.1. The

electrokinetic potential and contact angle measurements were used to show the

adsorption mechanism of sodium dodecyl sulfate and dodecylamine chloride on the

alumina surface. From these results, it was found that, sodium dodecyl sulfate and

dodecylamine chloride converted the alumina surface to a hydrophobic one at a pH less

than ZPC and a pH more than ZPC, respectively [13]. Zhenghe et al. [5] have used the

reverse flotation technology to float alumina from alumino silicates using a cationic

collector. The purpose of the presented work is to upgrade the alumina in gibbsite ore

to be suitable for industerial applications by using of sodium oleate as a flotation

collector. These industerial applications include aluminous chemicals, abrasives,

building stones, alumina refractories and aluminium extraction [4].

2. MINERALOGY

Based on petrography and EDX technique, several minerals were identified in the

studied samples as follows:

Gibbsite is turbid cryptocrystalline heavily stained black, orange and brown

colors. It occurs as radial aggregate embedded in marly matrix. The matrix is mainly

composed of equal amounts of fine-grained quartz and carbonate. It contains minute

grains of Fe and Mn oxy-hydroxides. The BSE of gibbsite is shown in Fig. 1.

Quartz is monocrystalline, medium-grained, well sorted and subround to well

round. The cryptocrystalline quartz occurs as filling pore spaces between grains and as

over growths. The BSE of quartz is shown in Fig. 2.

Kaolin occurs as amorphous dusts and patches with aggregate polarization in

crossed nicols. It mostly resulted from acidic ground water, whereas the unstable

minerals dissolve releasing Si, Al to the pore waters that can be produced by flushing

of the gibbsite by fresh water. The BSE of kaolin is shown in Fig. 3.

Halite appears as disseminated small cubic crystals encountered in gibbsite.

Calcite occurs as rhombohedral crystals, anhedral grains and aggregates radiating

fibrous. It exhibits twinning and perfect rhombohedral cleavages.

Hematite occurs either as disseminated anhedral crystals or occasionally as

minutes. It is steel-black with metallic luster in reflected light with a tendency to a

marginal red. Generally, the iron associated with manganese forming Mn-Fe minerals.

M. M. Ahmed et al.

446

Fig. 1: The BSE image of gibbsite

Fig. 2: The BSE image of quartz

Fig. 3: The BSE image of kaolin


Gypsum exists as white massive and acicular prismatic crystals of different

shape and sizes. They exhibit aggregates of subhedral to anhedral crystals having

vitreous luster. Few crystals show the perfect cleavage of the growing crystals. The

occurrence of gypsum in the studied gibbsite can be explained as due to upward

movement of brine by capillary action and its evaporation at the surface.

3. EXPERIMENTAL WORK

3.1 Materials

The samples used in the present work were obtained from the Southwest of Sinai (Abu

Thur). A batch representative sample of about 250 kg was taken from the gibbsite

zone. The obtained sample was split to smaller ones and crushed to minus 10 mm in a

laboratory jaw crusher. The crusher product was sieved on a screen of 200 µm size.

The screen oversize (+200 µm) was ground in a closed ball mill to minus 200 µm.

Grinding conditions were as follows: slope of mill = 0°, ball-to-ore ratio = 3:1, mill

speed = 60 rpm, and grinding time = 20 minutes. The crushing and grinding flowsheet

of the head sample is illustrated in Fig. 4.

A representative sample was taken and chemically analyzed for the

determination of the different constituents. The chemical analysis of the studied head

sample is given in Table 1

3.2 Reagents

All the flotation tests were carried out using sodium oleate as a collector. Sodium

hydroxide was used as a pH modifier. Tap water was used for maintaining the

flotation pulp at a constant level, as well as, all the other experimental purposes.

3.3 Apparatus

Laboratory flotation tests were carried out in a 3000 ml Wemco Fagergern cell. A

hand skimming was used to skim the froth.

3.4 Experimental Procedures

The impeller speed, during all flotation tests, was fixed at 1250 rpm and an aeration

rate of 6 L/min was used. The total conditioning time was 15 min. The studied

operating parameters were the pH value, collector dosage, pulp density and feed

particle size. The gibbsite sample was added slowly and then conditioned with water

for 5 minutes.

The pH value was adjusted at the end of the initial conditioning period and

allowed to condition for 5 minutes with the pulp. The collector dosage was added at

the end of the second conditioning period and allowed to condition for 5 min with the

pulp prior the aeration. The air supply valve was gradually opened. The required pulp

level was maintained constant during the flotation experiments.

In each experiment, and after allowing 15 sec for the froth to form, one

concentrate was collected during batch flotation tests. The collected products

M. M. Ahmed et al.

448

(concentrate and tailings) were dried, weighed and chemically analyzed. The

component recovery of alumina and other constituents in the concentrate and tailings

were calculated.

R.O.M

Gibbsite ore

Representative sample

+200 µm

200 µm screen

Jaw crusher

set at 10 mm

200 µm screen

Ball mill

+200 µm

-200 µm -200 µm -200 µm

-200 µm

To flotation process

Fig. 4: Crushing & grinding flow-sheet of the head sample

Table 1: Chemical analysis of the head sample of gibbsite ore from South

Western Sinai (Abuthur)

Al2O3% SiO2% Fe2O3% CaO% P2O5% MnO% K2O% Na2O% Traces% L.O.I.% Total%

18.98 15.45 12.83 14.87 5.74 5.34 0.86 0.76 1.50 23.65 99.98

Sampling


4. RESULTS AND DISCUSSION

4.1 Calculations of Experimental Mass and Component Recovery of Flotation products

Using the mass percent and assays of alumina in feed, concentrate, and tailings, the

experimental value of the mass recovery and component recovery of alumina in

concentrate and tailings can be calculated as follows:

(1)

(2)

(3)

(4)

4.2 Results of Using Sodium oleate as a Collector

4.2.1 Effect of pH value

Table 2 and Figure 5a show the effect of pH value on the grade, component recovery

of alumina in the concentrate, and the mass recovery of concentrate. These

experiments were carried out at a particle size of –200 µm, pulp density of 150 g/l, and

collector dosage of 1500 g/t. From this figure, it is noticed that the best selectivity is

achieved at a pH value of 11, where the component recovery of alumina in concentrate

was 78.3%, the alumina grade was 40.1%, and the mass recovery of concentrate was

about 37.1%. The grade of alumina decreases from 38.1% at a pH value of 6 to 33.5%

at a pH value of 9, then increases at higher pH value. The zero point charge of gibbsite is at pH values of (8.1–9). Below these

values, the gibbsite surface is positively charged, thus, negatively charged ions of any

anionic collector may be potentially adsorbed in this pH range [13]. Above the zero

point charge value, the surface of gibbsite is negatively charged, so ions of any anionic

collector are repelled from the gibbsite surface. Sodium oleate (C17H33COONa) is an

alkali soap of oleic acid. Although sodium oleate is an anionic collector, it floats

gibbsite above the isoelectric point of gibbsite. This indicates that, chemisorption

rather than the electrostatic interaction occurs between the oleate anion and gibbsite

surface [14, 15].

The isoelectric point of alumina shifts in the presence of sodium oleate and this

shift of the isoelectric point of gibbsite is attributed to the chemisorption of sodium

oleate on the gibbsite surface. It can be noticed that the component recovery of

alumina decreases from 73.9% at a pH value of 6 to 57.4 % at a pH value of 8, then

increases at higher pH values.

F

C100. (c)R econcentrat ofrecovery Mass

m

F

T100. (t)R tailingsofrecovery Mass

m

F.f

C.c100. (c)R econcentratin recovery Component

c

F.f

T.t100. (t)R sin tailingrecovery Component

c

M. M. Ahmed et al.

450

Table 2: The effect of pH value on the mass recovery, grade, and component

recovery of alumina

20

30

40

50

60

70

80

5 6 7 8 9 10 11 12

pH value

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %

mass recoverygradecomp onent recovery

Fig. 5a: Effect of pH value on the grade, component recovery of alumina in the

concentrate, and the mass recovery of concentrate.

Figure 5b shows the effect of pH value on the grade, component recovery of

alumina in the tailings, and the mass recovery of tailings. From this figure, it is

Exp. No.

pH value Product Mass Alumina, %

Recovery, % Grade Recovery

1

6

Concentrate 36.9 38.1 73.9

Tailings 63.1 7.9 26.1

Feed 100 19 100

2

7


Tailings 65.8 10.4 36.2

Feed 100 19 100

3

8


Tailings 66.3 12.2 42.6

Feed 100 19 100

4

9


Tailings 66.3 11.6 40.7

Feed 100 19 100

5

10


Tailings 64.2 9.8 33.1

Feed 100 19 100

6

11


Tailings 62.9 6.6 21.7

Feed 100 19 100


revealed that the component recovery of alumina in the tailings increases from 26.1%

at a pH value of 6 to 42.6% at a pH value of 8. At pH values greater than 8, the

component recovery of alumina in the tailings is decreased. The grade of alumina

increases from 7.9% at a pH value of 6 to 12.2% at a pH value of 8, and then decreases

at higher pH values.

0

1 0

2 0

3 0

4 0

5 0

6 0

7 0

5 6 7 8 9 1 0 1 1 1 2

pH va lue

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %


Fig. 5b: Effect of pH value on the grade, component recovery of alumina in

tailings, and the mass recovery of tailings.

4.2.2 Effect of sodium oleate concentration

Table 3 and Figure 6a illustrate the effect of sodium oleate concentration on the grade,

component recovery of alumina in the concentrate, and the mass recovery of

concentrate at the optimum value of pH (11) obtained from the previous experiments.

These experiments were carried out at the same conditions of pulp density (150 g/l)

and particle size (–200 µm). From this figure, it is shown that the component recovery

of alumina increases from 61.8% at a concentration of 500 g/t sodium oleate to about

79.9% at a dosage of 2000 g/t, then decreases at higher collector dosages. The mass

recovery of concentrate is increased from 27.6% to 38.9% at the same concentrations,

and then decreases at collector dosages higher than 2000 g/t.

The grade of alumina decreases also from 42.5 % to 39.1% at the same

dosages, and then increases at collector dosages higher than 2000 g/t. There is a

gradual increase in the recovery of alumina with the increase of the collector

concentration up to a maximum value. This may be due to the rapid reaction, more

rapid approach of exchange adsorption equilibrium, and may also be attributed to the

powerful action of sodium oleate to produce a water–repulsion and monomolecular

layer on particle surfaces thereby imparting hydrophobicity to the particles [16,17,18].

An excessive addition of collector dosages leads to an inverse effect and hence

decreases the component recovery of alumina. This may be due to the development of

collector multilayer on the particles, reducing the proportion of hydrocarbon radicals

oriented into the bulk solution. The hydrophobicity of particles is reduced and tends to

float other minerals, reducing selectivity and floatability [17].

M. M. Ahmed et al.

452

Table 3 : The effect of sodium oleate concentration on the mass recovery, Grade,

and component recovery of alumina.

0

10

20

30

40

50

60

70

80

90

0 500 1000 1500 2000 2500 3000 3500

C ollector dosage, g/t

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %

m ass recoverygradecom ponent recovery

Fig 6a: Effect of sodium oleate dosage on the grade, component recovery of

alumina in concentrate, and the mass recovery of concentrate.

Table 3 and Figure 6b show the effect of sodium oleate concentration on the

grade, component recovery of alumina in tailings, and the mass recovery of tailings.

From this figure it is seen that, the component recovery of alumina in tailings decreases

from 38.2% at a concentration of 500 g/t of sodium oleate to 20.1% at 2000 g/t, then

Exp. No. Collector

Product Mass Alumina, %

dosage, g/t Recovery, % Grade Recovery

1

500


Tailings 72.4 9.9 38.2

Feed 100 19 100

2

1000


Tailings 66.3 7.8 27.1

Feed 100 19 100

3

1500


Tailings 62.9 6.6 21.7

Feed 100 19 100

4

2000


Tailings 61.1 6.3 20.1

Feed 100 19 100

5

2500


Tailings 64.1 4.5 24.5

Feed 100 19 100

6

3000


Tailings 65.6 4.9 26.3

Feed 100 19 100


increases at higher collector dosages. The mass recovery of tailings decreases from

about 72.4% to 61.1% at the same collector concentrations, and then increases at

collector dosages higher than 2000 g/t.

The grade of alumina in tailings is decreased from 9.9% at a concentration of

500 g/t to 4.5% at 2500 g/t, and then increases at higher collector dosages.

0

10

20

30

40

50

60

70

80

0 500 1000 1500 2000 2500 3000 3500

C ollector dosage, g/t

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %

mas s reco v eryg rad eco mp o n en t reco v ery

Fig 6b: Effect of sodium oleate dosage on the grade, component recovery of

alumina in tailings, and the mass recovery of tailings

4.2.3 Effect of pulp density

Table 4 and Figure 7a show the effect of pulp density on the grade, component

recovery of alumina in the concentrate, and the mass recovery of concentrate. These

experiments were carried out at a pH value of 11, 2000 g/t sodium oleate dosage, and

with the same above stated particle size (–200 µm). From this figure, it is seen that the

component recovery increases from 79.9% at a pulp density of 150 g/l to 95.9% at 350

g/l, then decreases to 94.9% at a pulp density of 400 g/l. The mass recovery of the

concentrate is increased from 38.9% to 66.8% at the same pulp densities, and then

decreased to 59.9% at 400 g/l. The grade of alumina in concentrate decreased from

39.1% to 27.3% at the same pulp densities, and then increased to 30.1% at 400 g/l.

Wills [17] has reported that as a matter of economics, flotation separation must

be carried out in as dense a pulp as possible with good selectivity and operating

conditions. The denser the pulp, the lesser cell volume required, since the

effectiveness of most reagents is a function of their concentrations in solution. The

pulp must be diluted enough to permit particle re–arrangement to proceed freely. Over

dilution should be avoided as it results in greater amount of water consumption,

reagent consumption and more equipment for each tone of ore treated [24]. Therefore,

250 g/l is chosen as the optimum value of the pulp density.

M. M. Ahmed et al.

454

Table 4: The effect of pulp density on the mass recovery, grade, and component

recovery of alumina

0

10

20

30

40

50

60

70

80

90

100

100 150 200 250 300 350 400 450

pulp density, g/l

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %


Fig. 7a: Effect of pulp density on the grade, component recovery of alumina in


Exp. No.

Pulp density, g/l



1

150


Tailings 61.1 6.3 20.1

Feed 100 19 100

2

200


Tailings 57.9 4.8 14.5

Feed 100 19 100

3

250


Tailings 55.2 3.3 9.9

Feed 100 19 100

4

300


Tailings 43.9 2.9 6.9

Feed 100 19 100

5

350


Tailings 33.2 2.3 4.1

Feed 100 19 100

6

400


Tailings 40.1 2.4 5.1

Feed 100 100 100


Figure 7b shows the effect of pulp density on the grade, component recovery

of alumina in tailings, and the mass recovery in the tailings. From this figure, it is seen

that the component recovery of alumina in tailings decreases from 20.1% at a pulp

density of 150 g/l to 4.1% at 350 g/l, then increases to 5.1% at 400 g/l. The mass

recovery of tailings decreases from 61.1% to 33.2%, and then increases to 40.1% at the

same pulp densities. The grade of alumina in tailings decreases from 6.3% to 2.3%,

and then increases to 2.4% at the previous pulp densities. The excessive pulp density

leads to a low level of mineral extraction in the concentrate and increase the recovery

of alumina in the tailings.

0

10

20

30

40

50

60

70

100 150 200 250 300 350 400 450

pulp dens ity g/l

Gra

de

or

Ma

ss o

r C

om

po

ne

nt

rec

ov

ery

, %

mas s reco v eryg rad eco mp o n en t reco v ery

Fig. 7b: Effect of pulp density on the grade, component recovery of alumina in


4.2.4 Effect of particle size

Table 5 and Figure 8a show the effect of particle size on the grade, component

recovery of alumina in the concentrate, and the mass recovery of concentrate. These

experiments were executed at a pH value of 11, 2000 g/t sodium oleate concentration,

and a pulp density of 250 g/l. From this figure, it is seen that, the component recovery

of alumina increases from 69.1% at a particle size of (–200+160) µm to 97.6 % at

(−80+63) µm, then decreases to 89.9% at (–63+0) µm. The mass recovery of concentrate increases from 22.4% to 64.7% at the same

particle sizes, and then decreases to 54.7% at (–63+0) µm. The grade of alumina in

concentrate decreases from 46.9% at a particle size of (–200+160) µm to 33.2% at

(−80+63) µm, and then increases to 34.5% at (–63+0) µm.

Particle liberation plays an important role in the flotation process, where

particles of various sizes do not float equally well. Increasing particle sizes may result

in longer induction times, a commensurate deterioration in floatability, as well as, it is

expected that the coarser particles would tend to settle in the lower part of the flotation

cell [19,20]. The recovery is maximum for some intermediate size ranges, with a

M. M. Ahmed et al.

456

distinct fall towards the extreme course and extreme fine ranges [16,21]. Vijayendra

[16] reported that, the particles should not be overground, as the recovery and

selectivity decrease markedly if the particles are finer than 5 to 10 microns.

Table 5 : Effect of particle size on the mass recovery, grade, and component

recovery of alumina.

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60 80 100 120 140 160 180 200

particle size, Mm

Gra

de

or

Mas

s o

r C

om

po

nen

t

reco

ver

y%

mass%graderecovery%

Fig. 8a: Effect of particle size on the grade, component recovery of alumina in


Exp. No.

Particle size, µm



1

-200+160


Tailings 77.6 6.1 30.9

Feed 100 15.2 100

2

-160+125


Tailings 70.2 2.5 14.5

Feed 100 15.5 100

3

-125+100


Tailings 59.4 1.8 7.1

Feed 100 18.2 100

4

-100+80


Tailings 48.5 1.7 3.9

Feed 100 21.5 100

5

-80+63


Tailings 35.3 1.5 2.4

Feed 100 22 100

6

-63+0


Tailings 45.3 4.7 10.1

Feed 100 21 100


Figure 8b shows the effect of particle size on the grade, component recovery of

alumina in tailings, and the mass recovery of tailings. From this figure, it is seen that, the component recovery of alumina decreases from 30.9% at a particle size of (–

200+160) µm to 2.4% at (–80+63) µm, then increases to 10.1% at (–63+0) µm. The

mass recovery of tailings decreases from 77.6% to 35.3% at the same particle sizes,

then increases to 45.3% at a particle size of (−63+0) µm. The grade of alumina in

tailings decreases from 6.1% at a particle size of (–200+160) µm to 1.5% at –80+63

µm, then increases to 4.7% at (−63+0) µm. It can be noticed also that the rise in

tailing assays in the coarser fractions might be read to denote simply a progressive

decrease in liberation and failure of coarse free mineral to float.

0

10

20

30

40

50

60

70

80

90

100

0 20 40 60 80 100 120 140 160 180 200

particle size, Mm

Gra

de o

r M

ass

or

Co

mp

on

en

t

reco

very

%

mass%

grade

recovery%

Fig. 8b: Effect of particle size on the grade, component recovery of alumina in


5. CONCLUSIONS

From the obtained results the following conclusions can be drawn:

1. Cationic collectors can not be applied for the gibbsite ore due to the presence

of significant amounts of quartz.

2. The optimum operating conditions were as follows: sodium oleate dosage =

2000 g/t, pulp density = 250 g/l, pH = 11 and a particle size of (– 100+80) µm.

3. At optimum operating variables of flotation process, a concentrate of 51.5%

mass recovery containing 40.1% alumina grade with a component recovery

96.1% was obtained. The assays of other ore constituents in the final

concentrate were as follows: 5.9% SiO2, 4.8% CaO, 5.4% Fe2O3, 3.2% P2O5,

3.5% MnO, 1.3% K2O, 1.2% Na2O, 2.7% traces, and 31.4% loss on ignition.

The component recoveries of SiO2, CaO, Fe2O3, P2O5, MnO, K2O, Na2O,

and traces in the final concentrate were 21.4%, 17.5%, 21.2%, 38.4%, 42.9%,

75.3%, 89.6%, and 86.9%, respectively. 4. The final product is suitable for many industrial purposes such as aluminous

chemicals (aluminum sulfate and sodium aluminate are used for water

M. M. Ahmed et al.

458

treatment and aluminum chloride is used in refining crude petroleum), abrasive

products (coated abrasives, sharpening stones and grinding wheels). The final

product may be also applied for alumina refractories and alumina extraction

after its processing by Bayer process.

REFERENCES

1. Helena, S. S., Teresa, W. C., Persio, S. S., and Perdo, K. K., "Thermal phase

sequences in gibbsite/kaolinite clay: electron microscopy studies", Ceramics

International, Vol. 31, pp. 1077-1084, 2005.

2. Burkin, A.R., "Production of aluminum and alumina", John & Sons, New York,

pp. 3–13, 1987.

3. O`Connor, D.J., ״Alumina extraction from non bauxitic materials״, Aluminum-

Verlag GmbH, Germany, pp. 1–10, 1988.

4. Bittencourt, L.R.M., Lin, C.L., and Miller, J.D., "Flotation recovery of high

purity gibbsite concentrates from a Brazilian bauxite ore", In: Advanced

Materials–Application Mineral and Metallurgical Processing Principles, Society

of Mining Engineers of AIM, pp. 77–85, 1990.

5. Zhenghe, X., Plitt, V., and Liu, Q., "Recent advances in reverse flotation of

diasporic ores–A Chinese experience", Minerals Engineering, Vol. 17, pp. 1007–

1015, 2004.

6. Hinds, S.A., Husain, K., and Liu, N., "Beneficiation of bauxite tailings", Light

Met., Vol. 54, pp. 17–30, 1985.

7. Andreev, P.I., Anishchenko, N.M., and Mishakenkova, N.P., "Mechanism of the

action of amines during the flotation of bauxite ore minerals", Tsvetnye Metally,

Vol. 18, pp. 13–17, 1975.

8. Bulut, G. and Yurtsever, C., "Flotation behaviour of bitlis kyanite ore", Int. J.

Miner. Process., Vol. 73, pp. 29– 36, 2004.

9. Doss, S.K., "Adsorption of dodecyltrimethylammonium chloride on alumina and

its relation to oil–water flotation", Min. Process. Extr. Metall., C195–C199,

1976.

10. Wang, Y., Hu, Y., He, P., and Gu, G., "Reverse flotation of silicates from

diasporic–bauxite", Minerals Engineering, Vol. 17, pp. 63–68, 2004.

11. Marie, C.J., Fabien, G., and Bernard, H., "Limitations of potentiometric studies

to determine the surface charge of gibbsite γ–Al(OH)3 particles", Journal of

Colloid and Interface Science, Vol. 6, pp. 1–11, 2005.

12. Murray, B.M., "Adsorption of fatty acid spin probes on amorphous alumina",

Journal of Colloid and Interface Science, Vol. 76, pp. 393–398, (1980).

13. Hu, Y. and Dai, J., "Hydrophobic aggregation of alumina in surfactant solution",

Minerals Engineering, Vol. 16, pp. 1167–1172, 2003.

14. Vamvuka, D. and Agridiotis, V., "The effect of chemical reagents on lignite

flotation", Int. J. Min. Process., Vol. 61, pp. 209-224, 2000.


15. Ye, H. and Matsuoka, I., "Oleate flotation of dickite from quartz with diluted

HCl preconditioning. I. Effect of HCl preconditioning", Int. J. Min. Process.,

Vol. 40, pp. 83-98, 1993.

16. Vijayendra, H.G., "Handbook on mineral dressing", 2nd ed., Vikas, New Delhi,

pp. 195-210, 1995.

17. Wills, B.A., "Mineral Processing technology", Elsevier Ltd., 7th ed., pp. 283-

293, 2006.

18. Ahmed, M.M., "Kinetics of Magara coal flotation", M.Sc. Thesis, Assuit

University, Egypt, pp. 73-78, 1995.

19. Feng, D. and Aldrich, C., "Effect of particle size on flotation performance of

complex sulfide ores", Minerals Engineering, Vol. 12, pp. 721-731, 1999.

20. Zheng, X., Franzidis, J.P., Jounson, N.W., and Manlaping, E.V., "Modeling of

entrainment in industrial flotation cells: the effect of solid suspension", Minerals

Engineering, Vol. 18, pp. 51-58, 2005.

21. Ucurum, M. and Bayat, O., "Effects of operating variables on modified flotation

parameters in the mineral separation", Separation Purification Technology, In

Press 2006.

المثلى لتعويم االلومينا من خام طفلة جنوب غرب سيناء الحاملة للحالةالوصول للجبسيت باستخدام

Sodium Oleate كمجمع

المتواجدة فيي ايام اللة ية اكمجمع لتعويم األلو مين Sodium oleate في هذا البحث تم دراسة تأثير ينيا فيي ايام اللة ية الحام ية ل جبسييت ل ح يو الحام ة ل جبسيت بجنوب غرب سيناء بهيد تركييا اولوم

ع ى منتج بموا ةات مناسبة ونتاج اولومينا عالية التركيا ومناسبة ل ناعات المات ةة.والتلبيقيات ال يناعيل المات ةيل. Bayerايادة تركيا اولومينا بها لكى تكيو مناسيبل ل يداو فيى عم يية

ونسيبة SiO2 = 15.45% ونسيبة Al2O3 = 18.98%أ نسيبة وقيد تبيي مي التح يي الكيميياعى ل عينيلFe2O3 = 12.83% و نسيبةCaO = 14.87% و نسيبةP2O5 = 5.74% و نسيبةMnO =

trace elements = 1.50%و نسيبة Na2O = 0.76%و نسيبة K2O = 0.86%و نسيبة 5.43% %23.65. = و نسبة فاقد الحرق

٬ 2000g/tجرعييية المجميييع = ٬( pH=11ا عنيييد وسيييل قاعيييد الظيييرو المث يييى تيييم الح يييو ع يهييي تييم الح ييو وعنييد هييذظ الظييرو 80µm+100-) وحجييم الحبيبييات = 250g/lكثافيية المع ييق =

= SiO2% و نسيبة 96.1 اسيترجا = بنسيبةAl2O3 = 40.1%ع يى ركياا بالموا يةات اوتييلب نسيبة

% ونسييبة 17.6بنسييبة اسييترجا = CaO = 4.8% % و نسييبة21.4 بنسييبة اسييترجا = 5.9%

Fe2O3 = 5.4% = و نسيبة 21.2بنسيبة اسيترجا %P2O5 = 3.5% = و 38.4 بنسيبة اسيترجا %

M. M. Ahmed et al.

460

بنسيييبة اسيييترجا = K2O=1.3%% و نسيييبة 42.9اسيييترجا = بنسيييبة MnO = 3.5% نسيييبة trace elements = 2.7% و نسيبة % 89.6 بنسيبة اسيترجا = Na2O = 1.2%% و نسيبة75.3

اسيييترجا ل كت يييل بالنسيييبل لكت ييية الايييام 51.5%وقيييد تيييم الح يييو ع يييى %. 86.9 بنسيييبة اسيييترجا = .المغذ لا ية التعويم

ويمكيي اسييتادام هييذا المنييتج فييي العديييد ميي األغييرال ال ييناعية مثيي المييواد الكيمياعييية المسييتادمة فيييي المسييتادمة فييي تكرييير المييواد و ula su dosu adim و aluminum sulfateمعالجيية المييياظ و كذلك ال نةرة و ناعة الحراريات و ناعة األلومنيوم بعد aluminum chlorideالبترو مث ا Bayerتركياها بلريقة

Documents

FLOTATION OF ALUMINA FROM GIBBSITE BEARING–SHALE OF … · 2014-07-03 · However, the flotation of gibbsite by using amines as collectors can not be applied for the south western